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By James W. Hendrix

A survey of endomycorrhizal fungi occurring on orphan coalmine lands in Western and Eastern Kentucky revealed a great diversity of species present. Thus a large number of fungi are available for isolation and evalution for enhancing revegetation of mined land. Endomycorrhizal fungi isolated from orphan mines stimulated growth of sweetgum seedlings on mine soil at low fertilisation but inhibited growth at high fertilization; this growth inhibition at high fertilization did not occur with plants growing on peat-perlite. Banding of endomycorrbizal inoculum was not advantageous, compared with incorporation of inoculum throughout the growth medium, for production of mycorrhizal sweetgum seedlings. Newly reclaimed mine spoils were nearly nonmycorrhizal, compared with land reclaimed four or eight years earlier. A procedure for isolation of single-spore cultures of endomycorrhizal fungi was developed; this procedure is necessary for studies of ecotypic variability. A practical procedure using container-grown Pisolithus tinctorius-infected shortleaf pine seedlings for production of large numbers of bareroot pine seedlings infected with P. tinctorius for mined land reclamation was demonstrated.

Jan 1, 1982

By N. Kuroda

A realistic simulation model has been used for designing longwall mining, especially for solving operational problems. This paper describes background, development, validation and some applications of the simulation model for mining people who are unfamiliar with O.R. or computer technique, as a simple planning tool. The following items were established: logical analysis and algorithmic design for practical longwall mining, structure of hypothesis behavioral organism, input parameters, using practical knowledge and operational statistics of the mechanized longwall system experienced for many years.

Jan 1, 1982

By R. H. Sprute

This report describes a feasibility study to consolidate coal waste sludge in a 100-acre pond by applying direct current between floating and buried electrodes. The study was conducted by the Bureau of Mines under a cooperative agreement with Washington Irrigation and Development Co., located near Centralia, Wash. Part 1 covers results of laboratory tests at the Bureau's Spokane Research Center and a small-scale field test at the cooperator's coal preparation plant in western Washington. The process is efficient and cost-effective when power is applied at low current density. A proposed system, fully described in part 2, has electrodes covering 21 acres of pond surface radially arranged around a central power platform. Densification of pond contents would require 2 to 3 years of treatment. Total cost of equipment, labor, and material is approximately $500,000.

Jan 1, 1982

By S. S. Wang

Dialkyldithiophosphinates are a new class of flotation collectors for base metal sulfide ores. Structurally, they are quite different from the conventional collectors such as mercaptans, xanthates, thionocarbamates and dithiophosphates. Collectors Structures [Mercaptans R-SHS Xanthates R-OCS-M+ Thionocarbamates R-OCNH-R Dithiophosphates 2 R-CCNH-R' Dithiophosphinates (R) 2-PS-M+ ] In the above structural formula, R and R' stand for a hydrocarbon radical, and M+ is an alkali metal or ammonium cation.

Jan 1, 1982

By David F. Gibson

This paper presents an outline of several models that have been developed to aid surface mine planners evaluate alternative mining methods and dragline sizes. The models presented belong to a larger group of models that are incorporated in the SEAMPLAN computerized mine and reclamation planning system. SEAMPLAN was developed in the College of Engineering at Montana State University. Initial funding for the development was derived from the USDA Forest Ser- vice and EPA. Enhancements to SEAMPLAN are currently being funded by DOE (Contract DE-AC22-80ET-14173). SEAMPLAN is a comprehensive surface mine and reclamation planning tool that contains a number of major modules: Data Entry and Review, Production Analysis, Reclamation Analysis, and Impact Analysis. The Production Analysis Module contains programs that enable the user to evaluate alternative methods of removing overburden in a surface coal mine and then to assess auxiliary activities required to support the overburden removal operation. A set of these programs addresses the use of draglines for removing overburden. Six basic dragline mining techniques have been modeled. For one-pass dragline operations four methods can be analyzed: (i) simple side cast, (ii) side bench, (iii) extended bench, and (iv) extended bench with side bench. Two methods of two-pass operations are modeled: (i) single dragline, and (ii) tandem draglines.

Jan 1, 1982

By Joel K. Waterland

GENERAL DESCRIPTION Open cut-and-fill stoping for many years was prob¬ably the most widely used mining method in under¬ground metal mines. Then for a time this method was largely supplanted by the blasthole stope. It again be¬came popular as many mines reached depths or grades where methods requiring large open voids to remain open for extended periods of time became unsuccessful, often as a result of excessive dilution. The open cut-and-fill method is very flexible and is readily adaptable to almost any ore body. The standard application requires that a slice of ore usually 2.4 to 3 m (8 to 10 ft) thick be removed from the back of the stope, and as the ore is taken down, the back is dressed and rockbolted. After the back is secured, the broken rock is removed through rock passes to the level below. When the rock has been removed, the rock passes are extended upward the height of the ore removed, the stope is backfilled, and another cycle is mined. This method is best employed in plunging ore bod¬ies with considerable vertical extent, ore areas that re¬quire selective mining, ore areas where weak wall con¬ditions exist, and ore bodies that have an ore value that will carry this relatively expensive mining method. Blast¬hole stoping, shrinkage stoping, and other mining meth¬ods that do not employ rock passes in a stope are not efficient in plunging or flatly dipping ore bodies because the footwall makes ore removal quite difficult. Since mining is accomplished by taking down slices of the back, only small areas of the wall rock are exposed at any one time, and these only for short periods. Due to limited back height, areas of uneconomic rock may be left in place, or they may be mined and the material gobbed in the stope. Because the miners in the stope must work under freshly blasted areas, the amount of ground control is usually great. The volume of rock that is broken during one section of mining is relatively small and the amount of nonproductive work required is high. This results in limited productivity for the scope and, be¬cause of the nonproductive work that must be done on a regular basis, the production from the stope may be quite cyclical. SUITABLE ORE BODIES The open cut-and-fill method may be adapted to al¬most any type of ore body with a relatively high vertical extent. The ore body must be accessible at both top and bottom as well as at regular intervals throughout its vertical extent. Although adaptable to most ore bodies, the method is probably best employed where the ore has poor con¬tinuity and where most types of bulk mining would pro¬duce excessive dilution. In areas of poor ore continuity, the capability of continuous and extensive sampling dur¬ing the mining of each cycle makes this method very desirable. This capability also minimizes the amount of evaluation sampling that must be done before mining is started. Because of the extractive system used, the size and shape of the stope may be as readily changed as the sampling mandates. Probably the only ore characteristic demanded is that the ore has strength enough to be sup¬ported through the use of rockbolts or cable bolts dur¬ing the mining and backfilling cycles. Good planning, systematic sampling, and careful supervision will pro¬duce a product with less dilution than any other open stoping method. PLANNING Evaluation Planning Once it has been decided that the open cut-and-fill method would be the most efficient for mining a par¬ticular ore body, the next considerations would probably be the availability of an economical backfill material and the selection of an efficient transport system for this material. Although hydraulically transported mill tail¬ings are the most widely used product, this is not always practical due to mill location or the quality of the tailings. In such cases, backfilling may be used. The type of backfill and the type of equipment used will determine if a floor or capping on the backfill is required to minimize dilution during ore removal. The early selection of rock removal equipment is im¬portant since equipment usually determines the amount of development work required to bring a stope into pro¬duction and the size of the openings needed. The size and continuity of the ore body will usually determine the type of loading equipment. The use of slushers or load¬haul-dump (LHD) equipment captive in the stopes will minimize the amount of development. If the ore con¬tinuity is such that a ramp system for extraction can be used, the cost of development will be increased but the flexibility of continuous mining will minimize the cycli¬cal nature of the production. The height of the mining section usually is deter¬mined by the strength of the wall rock and the amount of back bolting required. Once this has been decided, the appropriate drilling equipment can be chosen. The number and sizes of the rock passes employed depends upon the type and size of the extractive equipment and the type of backfill that is to be used. Since the miners must enter and leave the stope each shift, the level inter¬val is usually maintained at approximately 45 m (150 ft). Access from the level above into the stope must be main¬tained at all times. The employees perform all the work in the stope and adequate ventilation must be provided. Stope Planning Due to the flexibility of the method and the vari¬ability of the ore zones, layout is usually done on a stope basis. In areas where continuity is a problem, the size of the stope is usually determined by the boundary of the ore (with all of the ore within that boundary being removed). In areas of good continuity where ramps are to be used, the length of the stope may be determined by the length of time each of the cycles (preparation, back¬fill, mining, and ore extraction) requires. The ramp work is then laid out so that access to the various parts

Jan 1, 1982

By H. R. Peterson

The control of autogenous grinding circuits is very similar to the control of conventional grinding circuits. In both circuits the classifier must be operated to achieve the required degree of mineral liberation. Crude ore grade and grinding mill density are important in both circuits. The biggest difference between the control of the two circuits is the pronounced effect that crude ore changes have on autogenous circuits. The importance of crude grade, weight recovery, size and hardness is greatly magnified because the crude ore in the autogenous circuit is also supplying the grinding media. Crude ore sizing is a particularly critical factor in achieving efficient primary autogenous milling. Because of the above, computer control is advantageous for autogenous circuits. Changes in crude ore may necessitate rapid changes in the entire circuit control strategy, and only a computer can facilitate these changes. Also, the techniques of control are very different for conventional and for autogenous grinding. Set-point control is usually satisfactory in conventional circuits, but the multiplicity of varying conditions in an autogenous circuit usually require some type of computer optimizing search routine. Autogenous mills can be separated into two general types. In primary autogenous mills the media is obtained from the mill feed, and in pebble mills the sized pebble media is added separately from the material being ground. PRIMARY AUTOGENOUS MILL CIRCUIT CONTROL - FIGURE 1 Power Draw Primary autogenous mills are usually run at full motor loading so production will be at maximum. Therefore, power draw must be closely controlled; this is done by adjusting feed rate. The versatility of computer control is extremely valuable for power control, especially during starting and stopping. To bring the mill power up as fast as possible at

Jan 1, 1982

By Thomas L. Starks

Surface soil/spoil Samples (0-10 cm) from experimental test sites in western North Dakota were aseptically collected during August 1975 and each month (May through September) during 1076 and 1977. Selected samples were Cultured in a defined inorganic medium and after six weeks the algae were identified. Chlorophyll u content of the soil/spoil was used as a measure of the algal abundance. Soil physical-chemical properties, species variety and abundance were statistically analyzed. All test sites revealed distinct soil algal communities and demonstrated a successional trend as evidenced by significant increase, in species variety and bionmass over time. Some soil factors, calcium, copper, and phosphorus affected species variety and abundance at more than one situ. Others were important at specific sites; silicon and aluminum at the South Beulah test site; manganese sodium and zinc at the North Beulah test site No. 1 and boron, cadmium and magnesium at the North Beulah test site No. 2. Topsoiled areas had higher species varieties and abundances when compared with control areas, as did plots treated with amendments such as leonardite and fertilizer Generally.

Jan 1, 1982

By K. P. V. Lei

The Bureau of Mines as part of its research effort to minimize the generation of hydrometallurgical waste, investigated a sodium sulfide leaching-electrolysis procedure for treating hydrometallurgical residues for elemental sulfur and base-metals recovery. Research was conducted on a leaching residue from a lead concentrate. The following procedure was developed. 1. The elemental sulfur was dissolved as polysulfide with an Na2S solution. 2. The resultant metal sulfide bearing residue was oxidatively acid leached in a cationic membrane electrolysis cell; the metals were solu¬bilized at the anode, and a base was generated at the cathode. 3. The elemental sulfur was recovered from the polysulfide solution by neutralizing with the acid in the anolyte. Byproducts of the neutralization were H2S and a sodium salt solution. 4. Sodium sulfide was regenerated by reacting the H2S byproduct with the base in the catholyte. 5. Lead as PbCL2 was recovered from the anode residue by brine leaching, and the other metals were recovered from the sodium salt solution as metal sulfides. 6. The barren filtrate from the metal sulfide precipitation constituted the recycled electrolyte. Ninety-eight percent of the elemental sulfur, 98 pct of the lead, and up to 80 pct of other accessory metals were recovered from a residue obtained from the FeCl3-NaCl leaching of a lead concentrate.

Jan 1, 1982

By N. Malakondaiah

Several growth chamber and laboratory experiments were conducted to study the germination behavior of plant species on mine spoils and plant growth responses to nutrients on spoils and topsoils. Experiments were conducted to study the germination potential of some native and introduced species under various conditions. The seeds of green needlegrass, side oats grama, thickspike wheatgrass, crested wheatgrass, birdsfoot trefoil and emerald crownvetch were less tolerant, whereas those of alkali sacaton. durar hard fescue, slender wheatgrass, yellow sweetclover and alfalfa were most tolerant to the sodic mine spoil conditions. Addition of leonardite improved the germination potential of some of the poor genninators. Slender wheatgrass and yellow sweetclover were used as test plants to study the fertility status and productivity potential of stockpiled topsoils and spoils collected from Center. Glenharold and South Beulah of North Dakota. The topsoils were neither saline nor sodic but low in N and P. The spoils were neutral to alkaline with SAIL of 12-32, EC of 1.8-5.5 mmhos/cm, Olsen&apos;s P< µ4g/g, and total N <0.07%. Application of N and P together to the topsoils increased the yield of wheatgrass and sweetclover by 88% and 70% of their respective maximum yields, whereas near maximum yields were obtained for both the species by the same treatments to the spoils. The average yields on topsoils were higher than the yields on spoils. The concentrations of major nutrients were higher in plants grown on topsoil, whereas the opposite was true with the concentrations of Mn, Zn, and Cu. Growth chamber experiments were also conducted to study the growth responses of alfalfa, oats and slender wheat grass by the application of Mn and Zn to a sodic coal mine spoil which was previously used for pot culture experiments. Significant growth responses to Mn and Zn were observed in alfalfa but not in oats or slender wheatgrass. Fertilizing the spoil with Mn along with NPK increased the yield of alfalfa to 85% and with Mn + Zn to 99%. Growth responses and plant composition results of this study confirm that the application of major nutrients to spoils can accentuate Zn and Mn deficiencies in some plants.

Jan 1, 1982

By Steven W. Schmauch

A mineral survey by the U.S. Bureau of Mines and U.S. Geological Survey indicate the Blanco Mountain RARE II area has low potential. Lode deposits consist of small, mineralized quartz or quartz-gossan veins. The veins are sparse and widely separated and lead, silver and gold content is restricted in amount and concentration. No gold or other potentially economic minerals were found in a placer deposit. There is no evidence of industrial mineral, geothermal, or energy resource potential in the study area. INTRODUCTION A mineral resource survey of the Blanco Mountain RARE II area (No. 5059), Inyo and Mono Counties, California was conducted in July 1981. This report summarizes the results and interpretations of the Bureau of Mines investigation.

Jan 1, 1982

By R. L. Handy

The goal of this research was to adapt the highly successful rock borehole shear test for use in deep, small-diameter boreholes. Operating somewhat on the principle of a rock bolt, the borehole shear test measures both the expansion pressure and pulling force to give discrete measures of rock shearing strength as a function of the applied normal pressure. A new instrument was designed and built to fit into a 2-3/8 inch diameter hole. Unfortunately the design became so complicated, the operation is nut reliable. This is attributed to constraints of the small hole, and added requirements of an in-hole anchor to push against, and a rotating shear plate cleaner for repeated tests down-hole. A clue to the d: vice complexity is its length, which is over 10 feet. Furthermore it employs concentric cylinders and one hydraulic cylinder bored into the piston of another, or it would be even longer. Meanwhile drilling technology has advanced to greatly reduce costs of drilling 5- or 6-inch diameter holes. The downhole tester therefore should be redesigned and simplified for use in a larger diameter drill hole.

Jan 1, 1982

By L. C. Lang

INTRODUCTION The introduction of 165-mm (61/z-in.) holes to underground mining operations has made possible the application of Canadian Industries Ltd&apos;s (CIL) vertical crater retreat (VCR) mining method, illustrated in the accompanying sketches. This unique and revolutionary new application of spherical charge technology (see the Appendix), when applied to primary stopes and pillar recovery, eliminates raise boring, slot cutting, and dilu¬tion of ore by backfill; greatly improves fragmentation; reduces labor and time requirements; eliminates uphole drilling and blasting; and minimizes or completely elimi¬nates damages by blasts to the walls and retreating backs of the stope or pillar. If vertical large diameter holes are drilled on a designed pattern from a cut over a stope or pillar to bottom in the back of the undercut, and spherical charges of explosives are placed within these holes at the calculated optimum distance from the back of the undercut and detonated, a vertical thickness of ore will be blasted downwards into the previously mined area. Repeating this loading and blasting procedure, min¬ing of the stope or pillar retreats in the form of horizontal slices in a vertical upwards direction until the top sill is blasted and the mining of a stope or pillar is completed. The VCR method is also applicable to drop raises and has the potential to replace all other raising methods under most circumstances. PILLAR RECOVERY Levack Mine Inco Metals Co., Ontario Div., provided the first opportunity for the method in pillar recovery. The Levack mine&apos;s high grade ore pillar No. 4800 on the 975-m (3200-ft) level was used for the production¬scale experiment (Figs. 1 to 3). The pillar was about 49 m (160 ft) long, 6 m (20 ft) wide, and 20 to 26 m (65 to 85 ft) high. The mined area on both sides of the pillar was backfilled with a 1:30 cement:sand mixture. The pillar was removed in two phases. In phase 1, the standard uphole method was used to blast down the 18-m (60-ft) long section of waste from the bottom of the ore into the undercut. From the pillar&apos;s top sill, 165-mm (61/a-in.) holes were drilled downward into the pillar, and by measuring the depth of the holes, the results of the uphole blast were determined and roof line 1 was established. The bottom of each hole was plugged, then filled with sand to place the center of gravity of each spherical charge (loaded from the top sill) at a predetermined optimum distance from the horizontal free surface. The charges were then detonated. After detonation, both draw drifts were full of extremely well-broken material. The depth of each hole was measured again, and the plot of these depths re¬sulted in roof line 2. The same blasting procedure was repeated and the resultant new back elevation was marked by roof line 3. The poor results of the initial uphole blast at one location (notice the peak in area 1) appeared to influ¬ence the subsequent new backs. A third blast success¬fully evened the back, and resulted in roof line 4. An unblasted slab averaging 6.3 m (20.9 ft) thick remained below the pillar&apos;s top elevation as the final sill. In all three spherical charge blasts fragmentation of the blasted material was extremely fine. The backfill was fully exposed on both sides of the now-blasted pillar. The backfill remained undamaged and the ore was not diluted by sand. The remnants of all the 165-mm (61/2 -in.) holes remained clean and undamaged, and the holes had well-defined bottoms that could be easily measured and plugged. Each blast took down a 3.9 to 4.2-m (13 to 14-ft) thick horizontal slab of ore. Productivity was three times greater than that of the previously practiced cut-and-fill method. Since this was the first such experiment, blasting the remaining 6-m (20-ft) thick final slab was the sub¬ject of some deliberation. If the described method was repeated, we could have ended up with a 1.8-m (6-ft) thick sill unsafe to work on. It was therefore decided to blast the whole sill using two spherical charges prop¬erly placed in each hole, but with the application of

Jan 1, 1982

By A. W. Chan

"An interactive computer simulation model of conveyor systems used in oil sands mining has been developed to aid in the selection of mining equipment and the evaluation of operating policies. The materials handling links between supply points, such as bucket-wheel excavators and draglines, and demand points such as bins and extraction plants are modelled in such a way that the conveyor model can be interfaced with any combination of these equipment submodels. Detailed calculations of the power demand of each conveyor section based on operating characteristics and conveyor design parameters are included to permit the close monitoring of energy requirements. Various interactive features are incorporated so that an operator can not only monitor, but also control, the mining operations in the model and experiment with different operating strategies.IntroductionThe interactive simulation model of conveyor systems described in this paper is one of a number of detailed computer models being jointly developed by the National Research Council and Canstar Oil Sands Ltd. on the mining equipment and operations used in oil sands mining. Reference I presents an overview of this whole series of models, their interrelationships and the underlying motivation for their development.A network of conveyors constitutes one of the practical ways for transporting mined material in an oil sands plant. The materials handling links between supply points, such as bucket-wheel excavators, bucket-wheel reclaimers and bins, and demand points such as dumps, bins and extraction plants can be represented by a series of interconnected conveyor sections. For example, the fluctuations of the feed rate at the bucket-wheel excavator, the flow of ore material through the conveyor network and the arrival of this material at the extraction plant can be simulated."

Jan 1, 1982

By Lawrence K. Nordell

This paper reviews practices used In the selection and sizing of belt conveyors and stacker systems commonly used in crushing and grinding plant facilities. Historical and modern methods of sizing this equipment are presented. The new methods Include: 1) probability theory applications for establishing component reliability, 2) "Monte Carlo" simulation techniques used in evaluating alternative system designs and in establishing service factors and corresponding availability Indices of each system, 3) quantifying the system and component operating characteristics using computer assisted modeling tools simulating rigid body and elasto-mechanic behavior for starting and stopping large and/or complex belt conveyors.

Jan 1, 1982

By J. W. Howland

A demonstration of dozer blades, designed for recontouring area-mined spoil and shown effective at Kansas and East Texas surface mine sites, was conducted at a Northeast Arizona surface coal mine to determine their effectiveness in different climate and soil conditions. The specialized tools were: 48-ft angle blade (on side-by-side tractor), 24-ft grading bar, modified 24-ft "U" blade (on side-by-side tractor), and a 13-ft deep "U" blade. Interest in the increased dirt-handling capacity of a previously built 22-ft blade, resulted in the design of a new Centerflow Blade of good versatility, able to maintain increased flow in otherwise difficult-to-move spoil. The demonstration involved 263 acres and showed a reduction of 50% in reclamation costs and 20% in fuel requirements for this mine site over conventional equipment. This savings resulted from improved technology, trained operators, and rigorous application of supervisory and monitoring techniques. Procedures and results of the demonstration are presented and analyzed here, along with guidelines for application of findings.

Jan 1, 1982

By L. K. Marshall

The importance of high quality underground mine ventilation is readily apparent. The principles of primary and auxiliary mine ventilation, however, are sometimes less readily apparent. The first article in this two-part series discusses some general aspects of primary and auxiliary ventilation. Next month&apos;s article will detail considerations involved with auxiliary mine ventilation.

Jan 1, 1982

By F. J. Arriagada

Where gold is present in a finely disseminated form within sulfide particles, the porosity and surface area of the calcined material strongly influence leaching since they control the penetration of solution inside the gold-bearing grains. On the basis of a stability diagram for the Fe-O-S in the temperature range of 400 to 750aC, the expected sequence chemical transformations associated with the oxidation of pyrite is discus-sed. By using analytical techniques which include scanning electron microscopy, X-ray diffraction and BET gas adsorption, the chemical and structural changes experienced by auriferous pyrite-arsenopyrite concentrates and pure pyrite samples are studied under an oxidizing roasting atmosphere of com-position 80% N + 20% O2 in the temperature range of 450 to 750°C. The results indicate that as the roasting temperature increases, the surface area of the calcines decreases sharply after a maximum value attained at about 450-500°C. SEM micrographs show that the needle-shaped tips which characterize the surface of the oxide product become wider and the size of the pores between them is decreased when the roasting temperature is in-creased from 550 to 750°C. This effect suggests that the formation of a more "closed" hematite product at higher temperatures is due to sintering associated with a faster growth and crystallization of the oxide. A strong correlation is shown between increasing roasting temperature and increase in gold lost in cyanidation tailings, an effect which may be partly attributed to the loss in surface area of the calcines.

Jan 1, 1982

By R. M. Board

This paper will focus on measurement and control of four main aspects of crushing, namely horsepower, bowl level, mantle position and feed rate, and will discuss related sensors and control techniques, some of them designed specifically for crusher control, that have proven successful and reliable in many installations.

Jan 1, 1982

By John De La Vergne, Morri Medd

"This paper describes the sinking of a circular concrete ventilation shaft through 110feet of water-bearing overburden. The brine circulation freezing method was used to stabilize the walls during excavation. In order to take advantage of the cold winter weather, a custom-built heat exchanger (natural radiator) was installed in place of the standard mechanical refrigeration plant.IntroductionIn the late fall of 1979, Les Mines Selbaie awarded a contract to the firm of 1.S. Redpath Limited to sink a ventilation shaft on its property in northwestern Quebec. The shaft was to be sunk through 110 feet of water-bearing overburden and approximately 90 feet of rock to a final depth of 200 feet below surface. The water table was measured at 16 feet below the surface elevation.After examining the soils report and diamond drill core from the shaft area, Redpath came to the conclusion that freezing was the best way to stabilize the overburden for sinking. LocationThe mine site is 420 statute miles due north of Toronto, just east of the Ontario/Quebec border (Fig. 1). Access to the site is provided by a development road built by the Quebec government from the Matagami/lames Bay Highway. History of Ground Freezing In the Mining IndustryIn 1862, shortly after the development of mechanical refrigeration, a mine shaft was sunk through an artificially frozen water-bearing formation in the United Kingdom. The first application of ground freezing in North America was for a Michigan iron ore shaft in 1888. Since that time, the freezing method has been used to sink hundreds of shafts throughout the world. Its best known application was to sink sixteen shafts through the Blairmore Formation in Saskatchewan.The same method has been used in every case. It involves chilling a calcium chloride brine solution and circulating the brine in concentric pipes, spaced between 3 and 6 feet around the shaft perimeter and usually 4 to 5 feet from the limits of the proposed shaft excavation."

Jan 1, 1982